Method for platinum recovery from materials containing rhenium and platinum metals

ABSTRACT

Hydrometallurgical methods are described for the isolation and recovery of platinum from rhenium-containing materials, and more particularly, from superalloys containing rhenium, platinum, and other metals. In addition, apparatuses capable of carrying out the hydrometallurgical methods and the product streams generated from the methods and apparatuses are described.

RELATED APPLICATIONS

This application claims benefit to provisional U.S. Application No.62/246,524, filed Oct. 26, 2015, the contents of which are incorporatedherein in their entirety.

FIELD OF THE INVENTION

The present disclosure relates to hydrometallurgical methods for theisolation and recovery of platinum from rhenium-containing materials,and more particularly, from superalloys containing rhenium, platinum,and other metals. The disclosure also relates to apparatuses capable ofcarrying out the hydrometallurgical methods and the product streamsgenerated from the methods and apparatuses.

BACKGROUND OF THE INVENTION

The present disclosure relates to a hydrometallurgical method for therecovery of platinum and other precious metals from complex feedstockscontaining rhenium. One example of a material containing both platinumand rhenium is a class of materials called superalloys. Creep-resistantrhenium-containing nickel and cobalt base superalloys were developed toprovide high temperature performance in severe environments, such asthose encountered in gas turbine engines and in blades for gas turbinegenerators. Rhenium confers creep resistance, i.e., resistance toplastic deformation at high temperatures, and corrosion resistance onthe alloys.

As used herein, the term “superalloy” means an alloy, such as a nickeland/or cobalt base alloy, containing chromium and rhenium, andcomprising one or more elements selected from tungsten, tantalum,zirconium, hafnium, molybdenum, yttrium, niobium, vanadium aluminum andplatinum. A typical superalloy contains 0.5 to 7% by weight rhenium,along with a major proportion (50 to 60%) nickel and minor amounts ofone or more of cobalt, chromium, aluminum, molybdenum, tantalum andtungsten (e. g. 2 to 10% of each). Virgin superalloy compositionstypically do not include platinum, which may be present as anundesirable impurity that is limited by product specifications.

Investment cast turbine blades are typically made from directionallysolidified and single crystal superalloys (e.g., Mar M 247, and CMSX-4).Superalloy turbine blades may be provided with a thermal barrier coat(TBC) made from a refractory oxide, such as platinum aluminide (forchemical or oxidative resistance), yttria-stabilized zirconia, or amixture of refractory oxides. To achieve a greater temperature gradient(and ultimately higher performance and fuel economy) between the hotgases flowing over the surface of the turbine blade and the internalblade structure, it is also common to cast in place a network ofinternal cooling structures. These complex channels are produced duringthe investment casting process with the use of cores held in place withpinning wires made of platinum or other noble metals having sufficientdimensional stability and oxidation resistance at de-wax and castingtemperatures. Nonlimiting examples of casting temperatures range from1300-1600° C., and nonlimiting examples of de-waxing temperatures rangefrom 700-1000° C.

When the component (e.g., turbine blade or vane) has reached the usefulservice limit for flight hours and/or cycles (e.g., 6000 cycles for asingle crystal turbine blade), the component is removed from service,examined and either refurbished and put back into service or recycledinto revert material. Due to the expensive and exacting nature of suchhigh temperature superalloys, which often contain rare elements such asrhenium, hafnium, tungsten, and tantalum in significant proportions,superalloy scrap is frequently recycled with virgin metals to produce analloy billet. However, some of these materials cannot be recycled backinto the revert stream when their composition does not meet stringentspecifications set by the manufacturers (e.g., because they arecontaminated, in some instances with excessive platinum). Theseend-of-life scrap materials are more valuable for their intrinsiccomposition and are therefore recycled for elemental recovery, and inparticular, rhenium recovery.

In some instances, the platinum contained in and on superalloycomponents such as turbine blades, gates, sprues, rises, and the likemay have a higher value than the other elements in the scrap alloy. Thesource of the platinum present in the alloy may be the pinning wire usedin the casting process, which may result in an alloy containing a highamount of platinum. The amount of platinum in such alloys is not limitedand may be as much as, for example, 0.34 wt %, 0.5 wt %, 1 wt % orhigher of platinum. Platinum can also be introduced into the alloy viathe remelting of platinum-aluminide coated airfoils. Standard industrypractice is to abrasive shot peen the end-of-life airfoils with cut wireor other abrasive media to remove surface contamination. The presentinventors have discovered that such treatment essentially welds andfurther diffuses the platinum into the base nickel alloy matrix, aslocally high temperatures and pressures are formed at the collision siteon the article.

Rhenium is present in these alloys and is conventionally recovered fromthese alloys in one of two ways—either by high temperature roasting ofatomized superalloy in air or oxygen to remove rhenium as a volatileoxide, Re₂O₇, or by hydrometallurgical electrochemical and/or chemicaldigestion wherein the superalloy is completely or partially dissolved inan acidic aqueous solution depending upon the dissolution parameters. Inhigh temperature roasting processes, platinum, if present, reports tothe calcined base metal oxides and may be recovered through traditionalhydrometallurgical approaches. In electrochemical or chemical digestionmethods, aggressive conditions using highly acidic solutions andoxidizing agents may solubilize all or a part of the platinum containedin the alloy feedstock. As a consequence, platinum enters the rheniumrecovery stream.

The rhenium in these alloys is recovered by loading into organicsolvents or upon a variety of resins, most of which are weak- or strongbase functionalized. Due to the similarity in chemical propertiesbetween rhenium and platinum, all or a part of the platinum follows therhenium throughout the process. Platinum competes for loading with weakand strong base resins and decreases the effective sorption capacity ofion exchangers thereby contaminating the resin-eluted rheniumconcentrate or strip liquor with platinum.

Superalloys are typically recycled to recover rhenium from the alloycomposition, but conventional rhenium recycling processes do not attemptto separate and recover platinum present in minor amounts from thedesired rhenium component because prior to the present invention, it wasnot known how to do so.

To this end, U.S. Patent Application Publication No. 2010/0126673 toDasan et al. and U.S. Pat. No. 5,776,329 relate to the roasting processfor removing rhenium from an alloy. WO2014158043 to Stroganov relates toa hydrometallurgical approach to removing rhenium from an alloy. Otherconventional methods for removing rhenium from an alloy are disclosedin, for example, U.S. Pat. No. 8,956,582 to Feron et al.; U.S. PatentApplication Publication No. 2003/0136685 to Stoller et al.; U.S. PatentApplication Publication No. 2013/0078166 and U.S. Patent ApplicationPublication No. 2011/0229366 to Luderitz et al; and U.S. PatentApplication Publication No. 2009/0255372 to Olbrich. The contents ofeach of the above documents are incorporated herein by reference intheir entirety. The conventional processes in the art, such as thoseabove, suffer from the problem that they do not provide a method foreasily separating small amounts of platinum that may be presented in therecycled alloy, in a form that is easily assayed.

Because such conventional hydrometallurgical recycling processes do notseparate rhenium from any platinum that may be present in recycledsuperalloy compositions, rhenium pellets produced fromhydrometallurgical recovery, which are consumed in the aerospaceindustry, often contain significant amounts of platinum impurity. Apractical and commercially viable method for the extraction of platinumfrom rhenium-containing superalloys would be highly desirable.

SUMMARY OF THE INVENTION

At least the following items are disclosed herein.

Item 1 is a method of separating platinum and rhenium comprising thestep of digesting an alloy feed comprising platinum and rhenium, whereinthe alloy feed is not contacted with a composition comprising more than250 ppm halides.

Item 2 is the method of item 1, wherein a step of separating resultingsolids comprising platinum from resulting liquids comprising rheniumfollows the step of digesting.

Item 3 is the method of item 1, wherein during the step of digestion,the alloy feed is not contacted with a composition comprising more than100 ppm halides.

Item 4 is the method of item 3, wherein during the step of digestion,the alloy feed is not contacted with a composition comprising halides.

Item 5 is the method of item 1, wherein the step of digesting comprisescontacting the alloy feed with sulfuric acid and a halide-free oxidant.

Item 6 is the method of item 5, wherein the sulfuric acid is present inan amount of 50-1000 g/L.

Item 7 is the method of item 5, wherein the halide-free oxidant isselected from the group consisting of air, ozone, oxygen, peroxide,persulfate salts, and mixtures thereof.

Item 8 is the method of item 7, wherein the halide-free oxidant isselected from the group consisting of peroxide and persulfate salts, ormixtures thereof.

Item 9 is the method of item 2, further comprising a step of recoveringrhenium from the liquids comprising rhenium.

Item 10 is the method of item 2, wherein the step of separatingcomprises a formation of a filter cake comprising platinum, wherein aconcentration of platinum in the filter cake is at least 2 times higherthan a concentration of platinum in the alloy feed.

Item 11 is the method of item 10, wherein a concentration of rhenium inthe filter cake is at least 50% lower than a concentration of rhenium inthe alloy feed.

Item 12 is the method of item 10, further comprising a step ofrecovering platinum from the filter cake.

Item 13 is the method of item 12, wherein the step of recoveringplatinum from the filter cake comprises digesting the filter cake in anoxidizing and complexing milieu.

Item 14 is the method of item 2, further comprising a step of recoveringplatinum from the solids comprising platinum.

Item 15 is a method of separating platinum and rhenium comprising thesteps of contacting a liquid comprising platinum and rhenium and achelating ion exchange resin; and adsorbing the platinum onto thechelating ion exchange resin.

Item 16 is the method of item 15, further comprising, prior to the stepof contacting the liquids comprising platinum and rhenium and achelating ion exchange resin, the steps of: digesting an alloy feedcomprising platinum and rhenium in a complexing ligand comprisinghalides in an amount sufficient to complex the platinum present in thealloy feed; and separating resulting solids from the liquid comprisingplatinum and rhenium.

Item 17 is the method of item 15, wherein the chelating ion exchangeresin is functionalized with thiourea or thiouronium groups.

Item 18 is the method of item 15, wherein the pH of the liquidcomprising platinum and rhenium is below 5.

Item 19 is the method of item 16, wherein the complexing ligand isselected from the group consisting of hydrochloric acid, a chloridecontaining salt, bromine and bromide salts, or chlorine.

Item 20 is the method of item 16, wherein the step of digesting furthercomprises digesting the alloy feed comprising platinum and rhenium in anoxidant capable of oxidizing both platinum and rhenium from theirmetallic zero-valence state to platinum's fourth oxidation state andrhenium's heptavalent oxidation state.

Item 21 is the method of item 20, wherein the oxidant is selected fromthe group consisting of peroxides, nitric acid and its salts, chlorates,chlorine, HCl, sulfuric acid, sodium chlorite, acids, oxygen, air, andmixtures thereof.

Item 22 is the method of item 21, wherein the oxidant is chlorate or amixture of HCl and sulfuric acid.

Item 23 is the method of item 21, wherein the oxidant is sodium chloriteand an acid.

Item 24 is the method of item 16, wherein a concentration of rheniumand/or platinum in the resulting solids is from 50 to 100% lower than aconcentration of rhenium and/or platinum in the alloy feed.

Item 25 is the method of item 24, wherein a concentration of rheniumand/or platinum in the resulting solids is from 90 to 100% lower than aconcentration of rhenium and/or platinum in the alloy feed.

Item 26 is the method of item 25, wherein a concentration of rheniumand/or platinum in the resulting solids is from 99 to 100% lower than aconcentration of rhenium and/or platinum in the alloy feed.

Item 27 is the method of item 15, further comprising, prior to the stepof contacting a liquid comprising platinum and rhenium and a chelatingion exchange resin, a step of decreasing an oxidative reductionpotential value of the liquid comprising platinum and rhenium.

Item 28 is the method of item 27, wherein the step of decreasing anoxidative reduction potential value of the liquid comprising platinumand rhenium comprises adding reducing agents to the liquid comprisingplatinum and rhenium.

Item 29 is the method of item 28, wherein the reducing agents arecapable of reducing platinum (IV) to platinum (II) without affecting theoxidation state of rhenium in solution (Re(VII)).

Item 30 is the method of item 29, wherein the reducing agents comprisesulfites, sulfur dioxide gas, or N₂H₄ salts.

Item 31 is the method of item 29, wherein the reducing agents compriseanhydrous sulfur dioxide, or a salt thereof, ascorbic acid, ethanol, orhydroxylamine hydrochloride.

Item 32 is the method of item 27, wherein the step of decreasing anoxidative reduction potential value of the liquids comprising platinumand rhenium comprises treating the liquid comprising platinum andrhenium with sulfur dioxide gas.

Item 33 is the method of item 15, further comprising, following the stepof contacting the liquid comprising platinum and rhenium and a chelatingion exchange resin, obtaining a liquid stream comprising a higherrelative concentration of rhenium to platinum than a relativeconcentration of rhenium to platinum in the liquids comprising platinumand rhenium.

Item 34 is the method of item 15, wherein at least 60% of the platinumpresent in the liquid comprising platinum and rhenium are adsorbed ontothe chelating ion exchange resin.

Item 35 is the method of item 34, wherein at least 90% of the platinumpresent in the liquids comprising platinum and rhenium are adsorbed ontothe chelating ion exchange resin.

Item 36 is the method of item 35, wherein at least 99% of the platinumpresent in the liquids comprising platinum and rhenium are adsorbed ontothe chelating ion exchange resin.

Item 37 is the method of item 15, wherein less than 10% by weight ofrhenium is adsorbed onto the chelating ion exchange resin, relative tothe amount of platinum adsorbed onto the chelating ion exchange resin.

Item 38 is the method of item 37, wherein less than 1% by weight ofrhenium is adsorbed onto the chelating ion exchange resin, relative tothe amount of platinum adsorbed onto the chelating ion exchange resin.

Item 39 is the method of item 38, wherein less than 0.1% by weight ofrhenium is adsorbed onto the chelating ion exchange resin, relative tothe amount of platinum adsorbed onto the chelating ion exchange resin.

Item 40 is the method of item 15, further comprising a step ofrecovering the platinum adsorbed onto the chelating ion exchange resin.

Item 41 is a system that separates platinum and rhenium comprising analloy feed comprising platinum and rhenium undergoing digestion in acomposition comprising more than 250 ppm halides.

Item 42 is the system of item 41, wherein the composition comprises morethan 100 ppm halides.

Item 43 is the system of item 42, wherein the composition comprises nohalides.

Item 44 is the system of item 41, wherein the composition comprisessulfuric acid and a halide-free oxidant.

Item 45 is the system of item 44, wherein the composition comprises thesulfuric acid in an amount of 50-1000 g/L.

Item 46 is the system of item 44, wherein the halide-free oxidant isselected from the group consisting of air, ozone, oxygen, peroxide,persulfate salts, and mixtures thereof.

Item 47 is the system of item 46, wherein the halide-free oxidant isselected from the group consisting of peroxide and persulfate salts, ormixtures thereof.

Item 48 is the system of item 41, comprising a filter cake comprisingplatinum, wherein a concentration of platinum in the filter cake is atleast 2 times higher than a concentration of platinum in the alloy feed.

Item 49 is the system of item 48, wherein a concentration of rhenium inthe filter cake is at least 50% lower than a concentration of rhenium inthe alloy feed.

Item 50 is the system of item 48, further comprising platinum recoveredfrom the filter cake.

Item 51 is the system of item 41, further comprising platinum recoveredfrom the alloy feed after digestion.

Item 52 is a system that separates platinum and rhenium comprising aliquid comprising platinum and rhenium; a chelating ion exchange resin;and platinum from the liquid comprising platinum and rhenium adsorbedonto the chelating ion exchange resin.

Item 53 is the system of item 52, further comprising, prior to thechelating ion exchange resin: an alloy feed comprising platinum andrhenium undergoing digestion in a complexing ligand comprising halidesin an amount sufficient to complex the platinum present in the alloyfeed; and a resulting separate streams of solids and the liquidcomprising platinum and rhenium.

Item 54 is the system of item 52, wherein the chelating ion exchangeresin is functionalized with thiourea or thiouronium groups.

Item 55 is the system of item 52, wherein the pH of the liquidcomprising platinum and rhenium is below 5.

Item 56 is the system of item 53, wherein the complexing ligand isselected from the group consisting of hydrochloric acid, a chloridecontaining salt, bromine and bromide salts, or chlorine.

Item 57 is the system of item 53, wherein the alloy feed comprisingplatinum and rhenium undergoing digestion is also digested in an oxidantcapable of oxidizing both platinum and rhenium from their metalliczero-valence state to platinum's fourth oxidation state and rhenium'sheptavalent oxidation state.

Item 58 is the system of item 57, wherein the oxidant is selected fromthe group consisting of peroxides, nitric acid and its salts, chlorates,chlorine, HCl, sulfuric acid, sodium chlorite, acids, oxygen, air, andmixtures thereof.

Item 59 is the system of item 58, wherein the oxidant is chlorate or amixture of HCl and sulfuric acid.

Item 60 is the system of item 58, wherein the oxidant is sodium chloriteand an acid.

Item 61 is the system of item 53, wherein a concentration of rheniumand/or platinum in the resulting stream of solids is from 50 to 100%lower than a concentration of rhenium and/or platinum in the alloy feed.

Item 62 is the system of item 61, wherein a concentration of rheniumand/or platinum in the resulting stream of solids is from 90 to 100%lower than a concentration of rhenium and/or platinum in the alloy feed.

Item 63 is the system of item 62, wherein a concentration of rheniumand/or platinum in the resulting solids is from 99 to 100% lower than aconcentration of rhenium and/or platinum in the alloy feed.

Item 64 is the system of item 52, wherein the liquid comprising platinumand rhenium further comprises reducing agents.

Item 65 is the system of item 64, wherein the reducing agents arecapable of reducing platinum (IV) to platinum (II) without affecting theoxidation state of rhenium in solution (Re(VII)).

Item 66 is the system of item 64, wherein the reducing agents comprisesulfites, sulfur dioxide gas, or N₂H₄ salts.

Item 67 is the system of item 64, wherein the reducing agents compriseanhydrous sulfur dioxide, or a salt thereof, ascorbic acid, ethanol, orhydroxylamine hydrochloride.

Item 68 is the system of item 52, wherein the liquid comprising platinumand rhenium has been treated with sulfur dioxide gas.

Item 69 is the system of item 52, further comprising a liquid streamcomprising a higher relative concentration of rhenium to platinum than arelative concentration of rhenium to platinum in the liquids comprisingplatinum and rhenium, which is obtained after platinum from the liquidcomprising platinum and rhenium has been adsorbed onto the chelating ionexchange resin.

Item 70 is the system of item 52, wherein the chelating ion exchangeresin has adsorbed at least 60% of the platinum present in the liquidcomprising platinum and rhenium.

Item 71 is the system of item 70, wherein the chelating ion exchangeresin has adsorbed at least 90% of the platinum present in the liquidscomprising platinum and rhenium.

Item 72 is the system of item 71, wherein the chelating ion exchangeresin has adsorbed at least 99% of the platinum present in the liquidscomprising platinum and rhenium.

Item 73 is the system of item 52, wherein the chelating ion exchangeresin has adsorbed less than 10% by weight of rhenium relative to theamount of platinum adsorbed onto the chelating ion exchange resin.

Item 74 is the system of item 73, wherein the chelating ion exchangeresin has adsorbed less than 1% by weight of rhenium relative to theamount of platinum adsorbed onto the chelating ion exchange resin.

Item 75 is the system of item 74, wherein the chelating ion exchangeresin has adsorbed less than 0.1% by weight of rhenium relative to theamount of platinum adsorbed onto the chelating ion exchange resin.

Item 76 is the system of item 52, further comprising platinum recoveredfrom the platinum adsorbed onto the chelating ion exchange resin.

BRIEF DESCRIPTION OF THE DRAWINGS

FIG. 1 is a schematic illustration of a process according to anembodiment of the present disclosure, in which platinum is separatedfrom rhenium as part of an insoluble solids residue.

FIG. 2 is a schematic illustration of a process according to anotherembodiment of the present disclosure, in which platinum and rhenium areco-dissolved and separated from insoluble solids.

DETAILED DISCLOSURE OF THE INVENTION

The present disclosure is based on the discovery of an efficient andeffective method for selectively recovering platinum fromrhenium-containing materials. The method is able to efficiently recoverplatinum from a variety of rhenium-containing materials containing othermetals such as nickel, cobalt, chromium, tungsten, tantalum, zirconium,hafnium, molybdenum, yttrium, niobium, and aluminum.

As used herein, the term “platinum” may include platinum and otherelements of the platinum metal family.

A rhenium-containing material is any material that contains Rhenium(Re). This includes waste, residue, ore, ore concentrate, byproduct,processed, and/or unprocessed material. Rhenium-containing materialsinclude nickel, cobalt, and/or molybdenum-containing manufacturingsludge residues, wastes, and byproducts. These materials have a physicalconsistency of a powder, sand or sludge and are typically comprised ofmetal compounds, metal alloys, metal grinding polishing fines, etchantcompounds, and mixtures thereof. Rhenium containing materials alsoinclude granular filter media, fibrous filter media, abrasive grindingmaterial and plasma deposition overspray particles. In one aspect ofthis disclosure, the rhenium-containing material is a superalloy waste,sludge, byproduct, or residue resulting from the manufacturing and/orsubsequent repair of high-temperature industrial turbines, turbinecomponents, superconductor components, vacuum plasma metal depositionprocesses, and bimetallic reforming catalyst materials.

The present disclosure is illustrated herein with reference torhenium-containing superalloys, but the disclosed embodiments can beapplied more broadly to any rhenium-containing material that alsocontains platinum, alone or in combination with other metals.

The method according to the present disclosure permits the efficient andcommercially viable separation and recovery of platinum fromrhenium-containing materials, including superalloys in which theplatinum is present in a small amount relative to rhenium, e.g., in aweight ratio of Pt:Re from about 0.002 to about 0.5. However, theconcentrations of rhenium and platinum present in the materials, andtheir ratio to one another, is not particularly limited, and thepresently disclosed methods may be used to separate platinum fromrhenium-containing materials in which these materials are present ineither large or small quantities. In conventional rhenium recoverymethods, it is difficult to separate such small amounts of platinum frommaterials containing rhenium, in order to avoid contamination of thefinal product. The present disclosure enables the recovery ofsubstantially pure rhenium and substantially pure platinum from alloyscontaining a mixture of these and other metals.

The present method provides for the isolation of platinum from an alloycontaining platinum and rhenium, in which the platinum and/or rheniumthat is produced does not contain more than 10% by weight, morepreferably not more than 5% or 1% by weight, and most preferably notmore than 0.1% by weight of any other metal. In preferred embodimentsthe present method provides for the isolation of platinum from an alloycontaining platinum and rhenium, in which the platinum and/or rheniumthat is produced does not contain more than 10% by weight, morepreferably not more than 5% or 1% by weight, and most preferably notmore than 0.1% by weight of any other material.

The present disclosure also includes the systems and parts and subpartsof the systems that achieve the methods disclosed herein. The systemsherein may include both the alloys and processed alloys herein, as wellas the intermediate products disclosed herein. The systems may alsoinclude the physical and chemical apparatuses necessary to carry out themethods disclosed herein. Where disclosures of materials and/orapparatuses are made herein, it should be understood that suchdisclosures also disclose the corresponding systems.

The phrase “substantially pure” means that material such as platinum hasa purity of about 90-99.9% by weight. The term “substantially free” of amaterial such as platinum similarly means that a compound or compositioncontains no more than about 0.1-10% by weight of the material such asplatinum.

Preparation of Metallic Feed

In preferred embodiments of the present method, an alloy containingrhenium, platinum and other metals is provided in a finely dividedparticulate form having a particle size range of about 2 microns to 300microns, and preferably from about 20 microns to 200 microns. Theparticulate form is preferably atomized. In general, lower the averageparticle sizes, and higher relative surface areas provide for fasterdigestion. In other exemplary embodiments, an alloy containing rhenium,platinum and other metals can be provided as larger pieces, for example,having a length of from about 5 cm to about 13 cm and an averagecross-sectional dimension of about 1.3 cm. In such embodiments, thelarge pieces may have a volume of about 20 cm³ to about 100 cm³.

In a preferred embodiment, a feed containing one or more of nickel,cobalt, vanadium, aluminum, titanium, hafnium, yttria, zirconium,tantalum, tungsten, chromium, molybdenum, rhenium and platinum (such asCMSX-4 alloy from jet turbine scrap manufacturing) is granulated oratomized to produce a homogeneous feedstock having a higher surfacearea, thereby providing for superior digestion kinetics. In thispreferred embodiment, the granulated superalloy feed is present infinely divided particulate form, in which the average particlesize/diameter may be, for example, about 2 to about 300 microns, andmore preferably about 20 to about 200 microns. The powder may, forexample, have an average surface area of about 10-50 m²/g. The surfacearea is understood to be contingent on the morphology of the particle.As an example, spherical or spheroid particles will have a minimizedsurface area to volume ratio, whereas flakey particles will have highersurface area to volume ratios.

The material may be sampled and analyzed for platinum, rhenium, andother elements of interest by total dissolution and gravimetric andspectroscopic analyses (e.g., NiS fire assay, and ICP-OES or ICP-MS)familiar to those versed in the art. The manner of conducting thisanalysis is not particularly limited, and any method known in the artmay be used. The feed in this embodiment is not limited, and personsskilled in the art will appreciate that the present process may besuitably applied to diverse feeds that may be treated with theembodiments disclosed herein, including without limitation thosecontaining platinum and rhenium in any proportion, alone or incombination with other elements and compounds.

Digestion of the Metallic Feed

In hydrometallurgical processes according to the present disclosure, thefeed material containing rhenium and platinum is digested by subjectingthe feed to strongly oxidizing conditions in an aqueous acidic solution,under conditions that are chosen to partially or completely dissolveplatinum, rhenium, and other metals present in the material or alloy.

The term “digest” as used herein means to wash, extract, or perform achemical reaction to separate a soluble element or compound from aninsoluble or relatively less soluble material. The phrase “insolublematerial” means an element in free form or a compound that is incapableof dissolving in a particular solvent, or that dissolves to a limitedextent, e.g., less than about 5%, or which dissolves and thenreprecipitates in the particular solvent, such as tungsten, which maydissolve in certain H₂O₂ systems to form soluble tungsten compounds, butthose tungstic compounds then reprecipitate.

In the digestion step, the alloy granulate or atomizate may be chargedinto a reactor of suitable construction to withstand the temperature andcorrosive environments of the digestion and sufficient capacity tocontain the solid and liquid charge. A suitable reactor may befiberglass reinforced plastic, a glass lined reactor, or a tantalumlined reactor, permitting the digestion conditions to be regulated fromroom temperature and atmospheric pressure to temperatures at orexceeding the boiling point of the reagent mixture at pressures rangingfrom about 1 bar to about 10 bar. The selection of the dissolutionchemistry dictates the pressures and temperatures employed.

In an exemplary embodiment, when the feedstock is an alloy that isfinely divided into particles having, an average size/diameter range ofabout 2 microns to 300 microns, the atomized particles areadvantageously subjected to strongly oxidizing conditions in an aqueousacidic solution at temperatures of from about 30° C. to about 80° C.,preferably from about 60° C. to about 80° C. and for a time period,which may range from about 1 hour to about 12 hours. If the feedstockmaterial containing platinum and rhenium consists of larger pieces, suchas pieces having a volume of about 20 cm³ to about 100 cm³, thefeedstock is preferably subjected to strongly oxidizing acid conditionsat temperatures of from about 40° C. to about 80° C. for a time periodof from about 2 days to about 10 days.

The feedstock solids content in the oxidizing acid solution ispreferably from about 5% to about 20% by volume, and more preferablyfrom about 10% to about 15% by volume. More concentrated solutions arepreferable to the extent they do not interfere with downstreamprocessing.

In the digestion step of hydrometallurgical processes according to anembodiment, the pH of the oxidizing aqueous solution and the oxidationreduction potential (ORP) of the solution are selected to selectivelydissolve relatively more soluble components or elements of the feedstockin the solution, in order to separate these elements from relativelyless soluble elements, or elements that are insoluble in the oxidizingaqueous solution. When the feedstock is an atomized superalloycontaining platinum and rhenium, the pH of the oxidizing solution ispreferably from about pH<0 to pH 1.5, most preferably from about <pH 0to 0.5.

Embodiment 1

In a first embodiment, platinum is recovered in solid form while rheniumis dissolved in an oxidizing solution. In the first embodiment, theplatinum present in a material such as a superalloy may be selectivelyseparated from the rhenium such that the platinum remains in theinsoluble digestion residues comprised of refractory metal oxides, SiC,hydroxides, and polyacids (e.g., tungstic acid and molybdic acid,hafnium oxides, tantalum metal and pentoxide, titanium (IV) oxide,zirconium (IV) oxide) and the solvated rhenium is dissolved in theaqueous acidic base metal stream. Platinum may be present in theinsoluble residues in the form of metallic platinum. The aqueous basemetal stream typically contains transition metals such as Ni, Co, Cr, Alas their sulfate salts.

Referring to FIG. 1, an incoming feed material containing platinum isprovided. The incoming feed material may be homogenized by, for example,being induction melted, preferably under vacuum or an inert gas blanketso as to preserve the rhenium from volatilization. The incoming feedmaterial may be granulated or atomized, so as to increase the surfacearea, thereby potentially providing for faster dissolution and/ordigestion. The incoming feed material may be, for example, CMSX-4 or MarM 247.

Again referring to FIG. 1, the incoming feed material may be weighed,sampled, and assayed. In this step of the method, the feed material isrepresentatively sampled. The sample may be solubilized via totaldigestion in a fluoropolymer pressure bomb with microwave-assistedpressure digestion (for example, in a solution of 3:1:0.5 concentratedHF:concentrated HNO₃:concentrated HCl). In this step, platinum and anyother elements of interest may be analyzed by, for example, inductivelycoupled plasma optical emission spectroscopy or mass spectroscopy(ICP-OES or ICP-MS), or other known methods. In this step of the method,all elements in the feed material may be dissolved, including but notlimited to the platinum.

Referring to FIG. 1, the feed material is provided into a reactor fordigestion. Water containing no halides (e.g., Cl⁻, Br⁻, I⁻, or F⁻) isadded to the reactor in an amount sufficient to ensure the solubility ofthe sulfate salts subsequently produced. Next, portions of sulfuric acidor phosphoric acid and any suitable halide-free oxidant (e.g., hydrogenperoxide, or nitric acid) is added to begin the reaction. Suitableconcentrations in this step may include, for example, 50-1000 g/L of,e.g., H₂SO₄. Other halide-free oxidants include, without limitation, forexample, air, ozone, oxygen, peroxide, and persulfate salts, andmixtures thereof. To prevent the formation of hydrogen gas, the ORPshould be maintained at or above 100 mV, as discussed below. Further, tofully solubilize the rhenium into solution as Re(VII), the ORP should beover 600 mV, and preferably over +800 mV. After the initial exotherms ofdilution and reaction subside, heat is applied so as to maintain thetemperature at or within, preferably, 10° C. of the boiling point of theacid mixture until the Re(VII) plateaus and the acid-soluble materialsdissolve. Other suitable temperatures for this reaction include, forexample 25-150° C. By avoiding the inclusion of halides in this step,zero-valence elemental platinum remains in the filter cake. In certainembodiments, all of the platinum present in the filter cake iszero-valence elemental platinum. In this respect, if halides are used inthe step of digestion, such use should not preclude that at least someportion of the Pt remains in the filter cake. To this end, the amount ofhalides used in the step of digestion is preferably less than 250 ppm,more preferably less than 100 ppm, and most preferably none. In apreferred embodiment, the digestion process is substantially free orfree of halides. If halides are used in this step, it is understood thatthe amount of platinum recovered in the filter cake may be adverselyaffected.

The digestion step in the first embodiment is preferably performed in anacidic oxidizing solution without the addition of halides that maysolubilize or complex the platinum. Platinum will not solubilize withoutcomplexing and oxidizing environments. Halides, such as chloride,bromide, and iodide constitute complexing ligands capable of maintainingplatinum in aqueous solution as a di- or tetravalent salt or complex.

To preclude the formation of inflammable hydrogen gas, the ORP of thedissolution mixture is preferably maintained from about +500 mV to about1000 mV, and most preferably from about +800 mV to about 1000 mV,relative to Ag/AgCl.

As described above, sulfuric acid, in concentrations ranging from about50-1000 g/L may be used as the acid, and preferably hydrogen peroxide,oxygen, ozone, or another oxidant not providing a complexing halide(e.g., Cl⁻, Br⁻, I⁻) may be used. More or less reagent may be required.Preferably the method is carried out at a sulfuric acid concentration ofabout 4-6 mole/liter, at a ratio of the acid to the granulated alloy ofabout (2-10):1 by volume, and a temperature of about 60° C.

In an alternate aspect of the present embodiment, electrochemicaldigestion may be used in addition to or instead of acid digestion, ifcomplexing media are not present as discussed above.

Referring to FIG. 1, the slurry formed in the digestion step is thenseparated into solids and liquids (e.g., soluble and insoluble)portions. Within the slurry, the rhenium may form perrhenic acid (HReO₄)and nonrefractory base metals may form their acid sulfates (including,but not limited to Ni(SO₄)₂, Co(SO₄)₂, Al₂(SO₄)₃, and the like), each ofwhich dissolve in and are part of the aqueous (liquid) stream. Thatstream is separated from the insoluble platinum-containing digestionresidues via centrifugation, decantation, filter pressing or any othermeans of effecting a solid-liquid separation.

As shown in FIG. 1, the rhenium-containing stream or liquor, which ispreferably substantially platinum free, may then be passed through aconventional weak base resin column system or a solvent extractionsystem so as to selectively sorb and recover the rhenium by knownmethods. Known methods of recovering rhenium from such streams may beused, including but not limited to ion exchange, solvent extraction,precipitation, and the like. An example of the recovery of rhenium withPurolite A170 weak base-functionalized macroporous resin is as follows:1000 L of a filtered (<1 um suspended solids), acidic (pH 0.5) aqueoussulfuric acid solution of ORP (550 mV) containing nickel, cobalt,aluminum, chromium and rhenium at 900 ppm was passed through three ionexchange beds containing 1.5 cubic feet (1 cu ft nonswollen) of anionexchange resin Purolite A170. The aspect ratio of each column was 1:6and each was comprised of clear PVC pipe having each end configured suchthat the resin could be removed for sampling. The rhenium-containingsolution was circulated through the resin columns at the rate of 3 BV/h(or 127 liters/hour; bed size being 1.5 cubic feet). Molybdenum had beenpreviously removed from the aqueous stream prior to contacting the A170with Purolite S957 Strong Acid Cation Resin. All loading, rinsing, andelution were conducted at room temperature. The resin was washed with 3BV of 5% w/v aqueous sulfuric acid for one hour to remove base metals,then 3 BV of deionized water to remove acidity from the resin. Therhenium outflow from the lag (final) column was less than 5 ppmindicating that full loading/breakthrough of the columns had not yetbeen achieved. The resin was further loaded in subsequent runs andstripped with 3 mol/liter aqueous ammonia to produce a dilution solutionof ammonium perrhenate suitable for further workup.

Referring to FIG. 1, after being separated, a platinum-containinginsoluble filter cake is obtained. In addition to platinum, theinsoluble filter cake may also include, for example, W, Ta, Zr, Hf, Mo,and the like. A typical example of the composition of such a filter cakefrom the digestion of CMSX-4 is seen in Table 1. An advantage to formingan insoluble filter cake in this manner is that the separation resultsin the removal of a majority, e.g., 80-90% by weight, of thenon-precious or non-refractory metals (e.g., not Pt, Ta, W, and thelike) present in the initial process stream, and the resulting filtercake therefore has a relative concentration of Pt that is much higher,e.g., at least about 7 times, preferably about 7-10 times higher, thanthe initial process stream. The separation may remove 50-90% by weightof the non-precious metals (such as, e.g., Ni, Co, and Re), preferably60-95% by weight, more preferably 70-99% by weight, and most preferably80-99% by weight or all of the non-precious metals. The relativeconcentration of Pt in the resulting filter cake may be from about 2 toabout 12 higher than the initial process stream, and in particularembodiments may be at least about 5, at least about 7, or at least about9 times to about 12 times more concentrated. A leached filter cake maycontain from about 20 to about 100 ppm Pt.

The concentration of rhenium in the resulting filter cake may be fromabout 50 to about 100% lower than the initial process stream, and inparticular embodiments may be at least about 80, at least about 90, atleast about 95, at least about 99, or at least about 99.5 percent less.It is preferable that the filter cake is free from or substantially freefrom rhenium.

The filter cake may be then be processed to yield a solution of, forexample, chloroplatinic acid from which platinum or other platinum groupmaterials may be recovered. For example, the insoluble residues orfilter cake, preferably containing substantially all of the platinum,may be treated with oxidizing and complexing media under strongly acidicconditions (preferably at a pH of 1 or below) with nitric acid,chlorine, hydrogen peroxide, or any other oxidant capable of maintainingan ORP at or above about +750 mV, and preferably above about +1000 mV.The acid(s) used are not particularly limited as long as a sufficientamount of halide is present (preferably at least a stoichiometricequivalent) to complex the platinum (or other platinum group metal(PGM)) and draw it into solution as a halegenoplatinate anionic complexof the form [PtX₆]²⁻ or, if in the bivalent state [PtX₄]²⁻ where Xconstitutes a halogen (e.g., Cl, Br, or I). The substantiallyrhenium-free and/or base metal free digestion slurry containing platinummay be separated from insoluble solids and the platinum is recovered viacomplex salt precipitation, ion exchange, electrowinning, cementation,or the like. Known methods of digestion may also be used in this step.The duration of the digestion can be varied, depending upon the alloy,its specific surface area, and the concentration of reagents andtemperature and pressure employed.

In another aspect, the insoluble residue so obtained may be subjected toa second digestion with aqua regia or other oxidizing and complexingmilieu that solubilizes, in part, or preferably in whole, thepartitioned platinum remaining. This platinum-containing solution isthen filtered and the platinum or platinum group metal recovered viaprecipitation, solvent extraction, ion exchange, to name but a fewmethods.

It is noted that the specific processing conditions of FIG. 1 areillustrative only, and do not limit the scope of the disclosure of FIG.1.

Embodiment 2

In a second embodiment of the disclosed process, platinum is recoveredfollowing codissolution with rhenium. In this embodiment the digestionmay be affected with a mixture of acids and oxidants and complexingagents (e.g., halides) to fully solubilize the platinum and rhenium,along with nonrefractory transition metals such as Ni, Co, Cr, and V astheir sulfate/chloride salts and complexes.

Referring to FIG. 2, in the second embodiment, the incoming feedmaterial and the step of weighing, sampling, and assaying the feedmaterial may be carried out as described above with respect to the firstembodiment.

With reference to the step of digestion shown in FIG. 2, in the secondembodiment both platinum and rhenium are dissolved. This is accomplishedby charging the feedstock into a reactor that is physically andchemically compatible with the reaction milieu, subject to constraintssimilar to the first embodiment. In this embodiment, a complexing ligandin quantity sufficient to complex the platinum contained in the solution(e.g., at least about a 10% excess of halide so as to produce [PtCl₆]²⁻)is used in the acidic digestion matrix. Preferred reagents arehydrochloric acid, a chloride containing salt, bromine and bromidesalts, or chlorine, or mixtures thereof. If digestion is conducted atatmospheric pressure, it is preferable to add the feedstock to an excessof the reagent; if digestion is conducted under pressure, it ispreferable to add an oxidant, discussed below, to the feedstock and acidmixture. In such a fashion, the rate of reaction and other parameterssuch as ORP, pH, and temperature may be more easily controlled.

During digestion, an oxidant capable of oxidizing both platinum andrhenium from their metallic zero-valence state to their fourth (Pt) andheptavalent (Re) oxidation state, respectively, may be used. Suitableoxidants for use in the digestion step include, for example, peroxides,nitric acid and its salts, chlorates, chlorine, oxygen, air, or anyother conventional oxidizer capable of maintaining the ORP at or abovethe prescribed level for full solubilization of platinum, or mixturesthereof. In a preferred embodiment sodium chlorate is the oxidant, whichoxidizes the chloride in HCl to chlorine, which solubilizes the Pt andthe Re. Another preferred embodiment utilizes a mixture of HCl andsulfuric acid and an oxidant. Another preferred embodiment utilizessodium chlorite, an oxidant, and an acid.

The ORP of the solution may be maintained during digestion at about +700mV to about 1200 mV, preferably from about +800 mV to about 1100 mV, andmore preferably from about +900 mV to about 1000 mV to ensure completedigestion of the platinum from the feedstock.

The platinum content in the digestion solution may be monitored byICP-OES or other appropriate analytical method. Digestion may beterminated upon achieving a plateau in Pt(IV) concentration, or uponreaching a predetermined concentration. When a plateau in Pt(IV)concentration is reached, substantially all of the metallic platinumpresent in the feedstock has been converted to soluble Pt(IV) saltforms, such as, but not limited to chloroplatinic acid (IV) orchloroplatinous acid (II). It is understood that if the platinumconcentration is at a steady state, so too is the rhenium concentration(e.g., rhenium dissolves before and during the dissolution of platinum,not after). If the amount of platinum in solution corresponds to theamount of platinum measured by assay, it may be presumed that all of therhenium has also been brought into solution.

Referring to FIG. 2, the slurry formed in the digestion step is thenseparated into solids and liquids (e.g., soluble and insoluble)portions. The process liquid stream may be separated from the solidresidue using a suitable method for liquid/solid separation effective toproduce a liquid stream containing, e.g., substantially all of the basemetals and platinum and rhenium in their highest typical oxidationstates (e.g., Ni(II), Cr(VI), Co(II), Pt(IV), Re(VII), Mo(VI), Al(III))and a solid residue comprised of e.g., refractory metals and theirrespective oxides. The solids-free solution may then be sampled andassayed.

Referring to FIG. 2, a solids filter cake may be obtained from theinsoluble portion of the slurry. This portion of the slurry may include,for example, W, Ta, Hf, Zr, some or all of the Mo, and other elements.The filter cake may be dried and then subjected to further processingto, for example, isolate elements of interest. The concentration ofrhenium and/or platinum in the resulting filter cake may be from about50 to about 100% lower than the initial process stream, and inparticular embodiments may be at least about 80, at least about 90, atleast about 95, at least about 99, or at least about 99.5 percent less.It is preferable that the filter cake is free from or substantially freefrom rhenium and platinum.

Referring to FIG. 2, solutions having high ORP values are potentiallydamaging and destructive to solvent extraction reagents, ion exchangeresins, and other downstream processes. To improve downstreamoperations, it is advantageous to perform, after solid/liquidseparation, an ORP adjustment step on the liquid stream. The liquidstream in this embodiment may include Re Pt, and, potentially, basemetal salts. Such a step may be carried out using as reducing agents,for example, sulfites, sulfur dioxide gas, or another reductant (e.g.,N₂H₄ salts) capable of reducing platinum (IV) to platinum (II) withoutaffecting the oxidation state of rhenium in solution (Re(VII)).Preferred reducing agents include anhydrous sulfur dioxide, or a saltthereof, and other reducing agents may also be used, including but notlimited to ascorbic acid, ethanol, hydroxylamine hydrochloride, and thelike. This step will lower the ORP of the solution, preferably to about700 mV or less, more preferably about 600 mV or less, even morepreferably to about 500 mV or less, and most preferably to about 380-480mV or less relative to Ag/AgCl to convert Pt(IV) to Pt(II) and protectdownstream processes from excessively oxidizing solutions, which maydamage, e.g., the chelating ion exchange resins. In a preferred aspectof the present disclosure, this action may be achieved by treating thesolution obtained with sulfur dioxide gas until the desired ORP isachieved. Such a reaction is quick and the byproduct is sulfuric acid.In addition, if nitric acid is present in the dissolution mixture, thetreatment step eliminates its presence. The reduction process may alsoserve to reduce hexavalent chromium to Cr(III).

The solution, having an ORP adjusted to this range, may then be sent toa platinum recovery circuit. FIG. 2 illustrates an embodiment in whichthe platinum is removed from the hydrometallurgical process stream priorto rhenium removal. The process stream containing metals including, forexample, platinum (II) and rhenium (preferably rhenium (VII)) is firstpassed through a chelating ion exchange resin, which may befunctionalized with thiourea or thiouronium groups. Examples of suchcommercially available resins include PUROLITE S920, PUROLITE S914,LANXESS LEWATITE MONOPLUS TP 214, and Dow Chemical® DOWEX 43600.00,among other suitable commercially available resins. Particularlypreferred resins are PUROLITE S920 or S914, or LANXESS LEWATITE MONOPLUSTP 214.

The solution, during the step of passing it through or contacting itwith a chelating ion exchange resin, preferably has a pH of 5 or belowin acidity, and the platinum loads to the resin preferentially overrhenium. The effect of the preferential loading of the platinum is a netconcentration increase of rhenium, based on the initial concentration ofthe rhenium in the acidic solution, while incurring only a minimalimpact on the rhenium units in the solution.

The chelating ion exchange resin may remove any amount of Pt, forexample, at least 10-100% by weight of the Pt present in the liquidstream, preferably at least 60-95% by weight, more preferably at least70-99% by weight, and most preferably at least 80-99% by weight or allof the Pt. It is preferable that the chelating ion exchange resinremoves less than about 10% by weight of rhenium, preferably less than5% by weight, more preferably less than 1% by weight, and mostpreferably less than 0.1% by weight of rhenium relative to the weight ofplatinum that is removed (e.g., adsorbed onto the chelating ion exchangeresin). Preferably, the chelating ion exchange resin is free orsubstantially free from adsorbed rhenium. Following exposure to thechelating ion exchange resin, the relative concentration of Re to Pt inthe resulting liquid stream may be increased proportionally to theamount of Pt removed.

In this embodiment, the platinum-rhenium containing digestion liquorscontaining acid-soluble base metals are first pumped through an ionexchange column containing the chelating agent so as to selectivelyremove platinum from the stream. The solution may be circulated throughthe ion exchange resin until a desired amount of platinum is recovered,preferably at least substantially all of the platinum in the solution.The amount of platinum present in the solution may be measuredspectroscopically from, e.g., the outflow of the final exit (lag)column. Flow and/or volume totalizers at the inlet and outlet of ionexchange columns may be used to determine the amount of platinum loadedby mass balance (see FIG. 2).

The ratio and concentration of the metal ions entering the ion exchangeresin, which may be present in a column, is not particularly limited.Platinum and rhenium having various ratios, including but not limited toratios of from 1:1 Pt:Re to 1:300 Pt:Re, and concentrations as low as,e.g., 500 ppb may be treated in this fashion, with time and processeconomics being a limiting factor.

In general, the ion-exchange resins are advantageously limited toneutral and acidic streams (having a pH about 7 or lower) sufficient toavoid degradation of the resin through hydrolysis of the thiourea orthiouronium moiety. The ORP of the process stream is preferably limitedto prevent oxidation of the thiourea or thiouronium moiety, and ispreferably not above about 800 mV in potential, more preferably notabove about 750 mV in potential, and most preferably not above about 600mV in potential.

It should be noted that in the present embodiment, acid-soluble basemetals refers to metallic elements other than those comprised of, e.g.,W, Ta, Hf, Zr, Nb, Ti, Ru, and to a lesser extent Mo, which is minimallysoluble under the described conditions. The flow volume, duration ofcirculation, and feed acidity may be varied according to the ionexchange column, and the fixed and mobile phases that are selected, asknown to persons skilled in the art. Process economics and the platinumlevel in the solution will direct the selection of specific variables.

Referring to FIG. 2, desorption or the incineration of the resin thenoccurs. The platinum-bearing chelating resin may be rinsed withhydrochloric or sulfuric acid (preferably about 4-6 M) to removeentrained base metals present in the resin void volume and weaklyabsorbed Re(VII). To recover the bound platinum, the resin may then beburned, with the possible loss of platinum and any bound rhenium units.

Preferably, the platinum is selectively eluted using conventionalmethods, for example, using a 1%-3% w/v HCl or H₂SO₄ solution containingfrom 1 g/L to the solubility limit of thiourea in such systems. Morepreferably, a 1% w/v HCl solution containing 50 g/L thiourea is used.The desorbate can then be treated with any suitable chemical reductant,electrowon onto cathodes, or hydrolyzed into platinum sulfide byprolonged hydrolysis of the thiourea in strongly alkaline solutions(e.g., at a pH of at least about 10 and at a temperature of from roomtemperature (e.g., about 20° C.) to boiling).

In other aspects of the disclosure, the eluent may be any eluent capableof complexing and removing the platinum from the resin in situ. If theplatinum is not removed in situ, the resin may be removed from thecolumns and incinerated for platinum value, as discussed above. Apreferred elution agent is 1-3% w/v HCl with 30-80 g/L thiourea. Otherelution agents that may be used include water-soluble thiocyanate salts,alkali or alkaline earth metal cyanides, or strong alkali bases.However, it is possible that strong bases and cyanides may destroy there-usability of the resin, and thus varying degrees of recovery may beachieved using such elution agents. It has been found that, generally, a10-12 bed volume (BV) elution with thiourea solution will achieveelution of about 95-98% of the bound platinum. A preferred temperatureof 60° C. may be used, but other temperatures may also be used,including temperatures from room temperature (e.g., about 20° C.) toabout 80° C.

Referring to FIG. 2, the platinum is recovered and refined following itsremoval from the chelating resin. The platinum produced in this manneris preferably substantially pure and free of rhenium and base metals andmay be refined to a commercially-acceptable purity specification usingmethods known in the art. Of particular import is that platinum soproduced is more easily refined than traditional hydrolytic methods,which fail to remove rhenium as a contaminant.

Referring to the embodiment shown in FIG. 2, platinum is recovered andrefined upstream, in part or in whole, prior to the rhenium recoveryprocess, such that rhenium units are not lost and rhenium recoveryprocesses, for example ion exchange and solvent extraction, are notcontaminated or damaged by the presence of excessive platinumconcentrations. According to this embodiment, at least about 80% of theplatinum present in the liquid stream may be removed prior to therhenium recovery. Preferably at least about 85%, more preferably atleast about 90%, even more preferably at least about 95%, and mostpreferably substantially all of the platinum in the liquid stream isremoved prior to the rhenium recovery process.

Referring to FIG. 2, after suitable platinum removal and decontaminationhas been achieved, the rhenium-containing process stream is processed torecover the rhenium. The rhenium-containing process stream willtypically contain base metal salts and Re(VII). A typical example of arhenium recovery cycle may include loading a weak or strong base resin(e.g., PUROLITE A170, A172) with rhenium, washing or otherwise elutingweakly bound species from the resin, rinsing off excess acidity, andthen eluting the rhenium with an alkaline solution capable ofregenerating the ion exchange resin and stripping/eluting the rhenium.The rhenium eluate thus obtained may then be processed further, viaevaporation and crystallization to ammonium perrhenate or othersparingly soluble perrhenate salts, so as to make salts, compounds, ormetallic rhenium powders of high purity for diverse applications.

In an aspect of the present disclosure, the step of rhenium recovery mayinclude the rhenium-containing stream being separately passed through anion exchange column for rhenium sorption and recovery. In such anaspect, an initial phase may include loading of the column with rheniumunits. A second phase may include washing or removing undesirablecontaminants, which may be bound weakly with low affinity, physisorbed,or else present in the void volume of the resin pores, from the columnby using, for example, acids having sufficient acidity to preventhydrolytic precipitation of base metals (e.g., 1-2 BV/h, for 1 hour),water (e.g., 1-3 BV/h, for 1 hour; or until the pH of outlet is fromabout 5 to about 7). A final phase may include elution via a 1-5 Mammonium hydroxide solution (e.g., 1-5 BV/h, for 1-2 hours or untilRe(VII) is eluted). The resin may then be finally rinsed free fromeluate using, e.g., distilled water, and may be reconditioned using,e.g., a dilute sulfuric acid wash (e.g., 3-5 BV/h, for 1 hour).

Suitable but not limiting examples of resins that may be used in therhenium recovery process are PUROLITE A170 or A172; both are weak basefunctionalized resins useful in rhenium recovery applications. Anothermore selective product for rhenium recovery and refining is based uponMolecular Recognition Technology, an example being SUPERLIG 188manufactured exclusively by IBC Technologies® of American Fork, Utah.

All commercially available rhenium sorption resins or solventextractants result in significant platinum uptake, since platinumco-loads into the resins or extractants alongside rhenium, as it doeswith all commercially available weak and strong base-functionalizedresins. Co-loading occurs because platinum and rhenium have similarchemistries, which makes them difficult to separatehydrometallurgically. The affinity of both Pt (II) and (IV) complexes inacidic conditions to weak base resins is comparable or greater thanrhenium (VII), making separation with that technology untenable. Asimilar situation arises with solvent extraction reagents based onamines (e.g., ALAMINE, ALIQUAT brands, and the like), alkyl phosphates(e.g., tributyl phosphate (TBP)), and alkylated phosphines and theiroxides (e.g., trioctylphosphine oxide (TOPO)); these extractants alsorecover rhenium in addition to and without distinction from platinum.For this reason, if platinum recovery is not substantially performedbefore rhenium sorption or solvation, the resin or solvent extractantagent is likely to become contaminated with platinum and requireextensive, expensive, and often environmentally undesirable separationmethodologies. According to the present disclosure, these disadvantagesof current recovery processes can be avoided, and accurate mass balanceand proper partitioning and separation of platinum from rhenium ispossible in streams processed via conventional hydrometallurgy.

Referring to FIG. 2, the rhenium eluate thus obtained is then processedfurther (e.g., by evaporation and crystallization to ammonium perrhenateor other sparingly soluble perrhenate salts) to make salts, compounds,or metallic rhenium powders of high purity for diverse applications. Asmentioned above, the insoluble residue left over from the digestion ofthe feedstock may contain Ta, Nb, Zr, Hf, Mo, and W—valuable elementsthat may be further recovered, separated, or otherwise purified.

The oxidative digestion of the described feedstocks can optionally beconducted using aqua regia (a mixture of nitric and hydrochloric acids)for superior digestion kinetics. In some exemplary embodiments, amixture of nitric acid and hydrochloric acid may be used. Thehydrochloric acid to nitric acid (v/v) in such exemplary embodiments maybe provided, on a concentrated acid basis, from about 1:3 to about 10:1.Furthermore, the hydrochloric acid, in some exemplary embodiments, mayconstitute a major portion of the acid mixture.

It is noted that the specific processing conditions of FIG. 2 areillustrative only, and do not limit the scope of the disclosure of FIG.2.

Example 1

A 100.0 g sample of superalloy granulate of undetermined mesh sizeproduced from the induction melting of platinum-aluminide-coated CMSX-4superalloy high pressure turbine blades was digested per the firstembodiment. The initial composition of the alloy is described inTable 1. The feedstock alloy was charged into a 2 liter round bottomPyrex flask equipped with an ORP meter/temperature probe and dosing pumpfor the addition of 35% w/v H₂O₂. 1.25 liters of 800 g/L sulfuric acid(with precautions taken to ensure the reagents were free of halides) wasadded into the reactor and hydrogen peroxide pumped in incrementally tomaintain an ORP at or above 500 mV relative to Ag/AgCl. No stirring wasemployed. After the initial exotherm subsided, the mixture was heated at80-90° C. for 35 hours. To check the completion of the reaction, thedosing pump was switched off and the reaction mixture examined visuallyto ensure no hydrogen gas was produced, which would be indicative ofincomplete digestion of the base metals. The pH of the solution remainedbelow 1 for the duration of the digestion and additional peroxide wasadded before filtration until an ORP of +800 mV was maintained. The darkgreen reaction mixture was then vacuum filtered through a pre-taredWhatman 42 filter paper to separate the insoluble residue. The residuewas rinsed with 5% w/v aqueous sulfuric acid until the filtrate wascolorless and then with deionized water until the filtrate was pH 6. Thefinal volume of the filtrate and all washing was 2.29 L. The dark greyfilter cake was then dried to constant mass (12.757 g), the filter paperremoved, and a 1 gram sample removed for microwave—assisted pressuredigestion in 25 mL of 3 parts 70% w/v HF to 2 parts 70% w/v HNO₃ to 1part 37% w/v HCl with 6 parts of deionized water. The filtrate from thesuperalloy digestion was dispensed into a pre-tared beaker and a 10 mLsample taken for density determination and analysis via ICP-MS. Theresults are summarized in Table 1.

The filtrate was found to contain substantially all of the soluble basemetals and rhenium but negligible platinum. The filtrate, having a pH of<1 was then passed through 50 grams of Purolite A170 macroporous weakbase resin loaded into a 30 cm×3 cm glass column with fitted disk torecover the rhenium. The resin had been preconditioned to the feed byequilibration for 5 h with 5 M H₂SO₄ solution. After the rhenium wassorbed onto the resin, the resin was washed with a 5% w/v aqueoussolution of sulfuric acid to rinse the base metals from the void volumeand then rinsed with deionized water to remove excess acidity. The resinwas then eluted with 10 bed volumes (approx. 1 L) of 3 M ammoniumhydroxide and the rhenium-containing eluate concentrated and thencrystallized to yield ammonium perrhenate. No platinum was found to bein the rhenium fraction so obtained.

TABLE 1 Composition of Cannon Muskegon Single Crystal 4 Superalloy Feed(CMSX-4) and streams derived therefrom Elemental composition (wt. %) NiCo Cr Mo W Hf Ta Al Ti Re Pt 62.00 8.75 6.41 0.61 6.00 0.11 6.50 5.601.00 2.71 0.31 After treatment per Embodiment 1 - insoluble residues(wt. %) — — 3.20 1.19 41.30 0.89 50.94 — — 0.04 2.43 % Recovery (wtpercent) <0.2 — — — — — — 0.2 99.98 After treatment per Embodiment 1 -aqueous Re stream (ppm) 27100 N.M. N.M. N.M. N.M. N.M. N.M. N.M. N.M.1170 0.2 Ammoniacal Re eluate (ppm) (before crystallization) <5 <5 <5 5245 N.M. N.M. N.M. N.M. 2657 NIL N.M. = not measured. Some recoveries arehigher than unity due to formation of oxides. NIL = No detectableamount, based on instrument used.

Example 2

Granulated superalloy material of the composition and form described inExample 1 was digested in a mixture of aqueous sulfuric acid (initially800 g/L) and hydrogen peroxide with the oxidation-reduction potentialmaintained at all times above 500 mV relative to Ag/AgCl. The reagentsused were checked for halide contamination with silver nitrate solutionto ensure no complexing halides were present. The digestion wasterminated after subsequent additions of hydrogen peroxide failed toincrease the Re(VII) concentration in the lixiviant (i.e., leachsolution) as measured by ICP-OES. The solution was filtered free ofsolids, the filter cake was washed well with dilute sulfuric acid (5%w/v) to remove base metals and soluble rhenium, and the rheniumrecovered with PUROLITE A170 resin. Elution of the resin with aqueousammonia produced an ammonium perrhenate stream substantially devoid ofplatinum (<1 mg/L), based on the detection limit of the measurementapparatus. The concentration of platinum remaining in the insolublefilter cake material was found to be nearly ten times higher than in thefeed material, corresponding well with the expected mass loss due tosulfuric-acid soluble elements. The filter cake was leached with sodiumchlorate and 6 M hydrochloric acid, filtered, and the filter cake rinsedwith dilute hydrochloric acid (3% w/v) until washings were negative tostannous chloride colorimetric test for platinum (an orange indicationon a test strip saturated with SnCl₂ solution). The filtrate wasweighed, its density determined, sampled, and analyzed via ICP-OES.Essentially all of the platinum (>99.5%) was removed from the insolublefilter cake. In this example, 682 kg of feed material containing 2114 g(3.1 g/kg) Pt was processed. The second step removed 2105 g of platinum.

Comparative Example 1

Granulated superalloy material was digested in a mixture of sulfuricacid, water and hydrochloric acid with hydrogen peroxide addition withthe oxidation-reduction potential maintained such that hydrogen gas wasnot formed during the dissolution. The feedstock contained 27 g/kgrhenium and 3.1 g/kg platinum after analysis of a sample and subsequentICP-MS determination. All of the rhenium and platinum reported to theaqueous reaction mixture as Re(VII) and Pt(IV), the content of themetals as reported in Table 1. The mixture was as separated frominsoluble solids (W, Ta, Zr, Hf, Ti oxides) by flocculation and filterpressing using 150 ppm nonionic polyacrylamide polymer and standardpressing conditions with two washes of the filter cake with distilledwater. The solids-free solution was circulated at ambient pressure andtemperature through PUROLITE A170 weak base resin to recover the rheniumuntil the rhenium outflow from the lag column was less than 1 ppm.Elution of the rhenium bound to the resin via 4 M aqueous ammoniaproduced a strip solution containing both platinum and rhenium, thus noseparation was achieved. Ammonium perrhenate produced from such asolution by evaporation and recrystallization was thereby contaminatedwith platinum.

Although the present disclosure has been made in detail with referenceto specific embodiments, materials and examples, it is not limitedthereby, and persons skilled in the art will be able to make variationsin the disclosed examples and description without departing from thespirit and scope of the disclosure as described in the claims.

It should be understood that the embodiments above are not limited tothe specific aspects disclosed above. In particular, aspects of each ofthe embodiments are available for use in other embodiments, and thespecific embodiments described herein should not be construed as beinglimited to only the conditions disclosed with respect to that specificembodiment unless otherwise noted.

We claim:
 1. A method of separating platinum and rhenium comprising thesteps of: digesting an alloy comprising platinum and rhenium therebyobtaining solids comprising platinum and liquids comprising rhenium,wherein the alloy is not contacted with a composition comprising morethan 250 ppm halides; and a step of separating the solids comprisingplatinum from the liquids comprising rhenium following the step ofdigesting.
 2. The method of claim 1, wherein a weight ratio of Pt:Re inthe alloy in the step of digesting is from about 0.002 to about 0.5. 3.The method of claim 2, further comprising a step of recovering rheniumfrom the liquids comprising rhenium.
 4. The method of claim 2, whereinthe step of separating comprises a formation of a filter cake comprisingplatinum, wherein a concentration of platinum in the filter cake is atleast 2 times higher than a concentration of platinum in the alloy. 5.The method of claim 4, wherein a concentration of rhenium in the filtercake is at least 50% lower than a concentration of rhenium in the alloy.6. The method of claim 4, further comprising a step of recoveringplatinum from the filter cake.
 7. The method of claim 6, wherein thestep of recovering platinum from the filter cake comprises digesting thefilter cake in an oxidizing and complexing milieu.
 8. The method ofclaim 2, further comprising a step of recovering platinum from thesolids comprising platinum.
 9. The method of claim 1, wherein during thestep of digestion, the alloy is not contacted with a compositioncomprising more than 100 ppm halides.
 10. The method of claim 9, whereinduring the step of digestion, the alloy is not contacted with acomposition comprising halides.
 11. The method of claim 1, wherein thestep of digesting comprises contacting the alloy with sulfuric acid anda halide-free oxidant.
 12. The method of claim 11, wherein the sulfuricacid is present in an amount of 50-1000 g/L.
 13. The method of claim 11,wherein the halide-free oxidant is selected from the group consisting ofair, ozone, oxygen, peroxide, persulfate salts, and mixtures thereof.14. The method of claim 13, wherein the halide-free oxidant is selectedfrom the group consisting of peroxide and persulfate salts, or mixturesthereof.
 15. A system that separates platinum and rhenium comprising: analloy comprising platinum and rhenium undergoing digestion in acomposition comprising not more than 250 ppm halides.
 16. The system ofclaim 15, wherein the composition comprises more than 100 ppm halides.17. The system of claim 16, wherein the composition comprises nohalides.
 18. The system of claim 15, wherein the composition comprisessulfuric acid and a halide-free oxidant.
 19. The system of claim 18,wherein the composition comprises the sulfuric acid in an amount of50-1000 g/L.
 20. The system of claim 18, wherein the halide-free oxidantis selected from the group consisting of air, ozone, oxygen, peroxide,persulfate salts, and mixtures thereof.
 21. The system of claim 20,wherein the halide-free oxidant is selected from the group consisting ofperoxide and persulfate salts, or mixtures thereof.
 22. The system ofclaim 15, comprising a filter cake comprising platinum, wherein aconcentration of platinum in the filter cake is at least 2 times higherthan a concentration of platinum in the alloy.
 23. The system of claim22, wherein a concentration of rhenium in the filter cake is at least50% lower than a concentration of rhenium in the alloy.
 24. The systemof claim 22, further comprising platinum recovered from the filter cake.25. The system of claim 24, further comprising platinum recovered fromthe alloy after digestion.